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By A. U. Seybolt, C. H. Mathewson

As is well known, many questions affecting the properties and uses of a metal cannot be answered without careful consideration of the state of purity realized in the various operations of preparation, refining, or special treatment that have brought the product to its final state. The occurrence of an impurity in minute quantity offers no guarantee that its effect will be inconsiderable because dissolved substances often modify the properties of a metal quite out of proportion to their concentration. In such cases the determination of solid solubility must be accepted as a problem of considerable difficulty, requiring the use of especially purified materials and methods not ordinarily used in the metallurgical laboratory. This aspect of the equilibrium between metals and their oxides is especially important and a search for fundamental data reveals little except conflicting opinion even concerning the systems of common occurrence and interest. Rhines and Mathewson1 in 1934 studied the solubility of oxygen in solid copper by annealing to equilibrium in air and gave reasons for the failure of other methods, notably the ordinary procedure of heat-treatment followed by microscopic examination, to yield reliable results. In this paper, a similar study has been made of the solubility of oxygen in cobalt. The simple theory of this method is outlined below.

Jan 1, 1935

By A. J. Fenn

Long experience with underground mining inspired the unusual idea of using an underground crusher-conveyor system as a substitute for surface haulage at the Cananea open-pit copper mine in Sonora, Mexico. Now implemented and working well, the system has born out the expectations of its originator, Cia. Minera de Cananea S.A. de C.V., that it would effect substantial cuts in the high cost of haulage. The Cananea deposit is a large, low-grade orebody above the mined-out Colorada pipe which was once one of the largest high-grade underground deposits in the district. Present facilities at the pit consist of: (1) A large open-pit mine capable of moving 80,000 tons of material per day with a shovel-truck, underground crusher and belt conveyor haulage system. (2) A 25,000-tpd concentrator equipped for four stages of crushing, rod mill-ball mill grinding with cyclone classification, and flotation in conventional machines. (3) A dump leaching process producing 1 million lb of copper per month in a tin-can scrap precipitation plant. (4) A two-reverberatory, six-converter smelter which can make 140 tons of blister copper per day. (5) A well-integrated contingent of supporting departments running the gauntlet from housing to power generation.

Jan 3, 1969

By R. G. Mallon, A. E. Lewis, R. L. Braun, C. J. Sisemore

Large ore-bodies, too deep to mine economically by underground methods, may be mined using nuclear explosives and a high pressure leaching technology. Breaking of a hypothetical ore-body with a series of nuclear explosives is described. The leaching would be accomplished by pumping oxygen into the flooded rubble under the normal hydrostatic pressure. The injection of oxygen gas into the bottom of the rubble chimney induces a circulation of the solutions which carries dissolved oxygen to all parts of the chimney. Copper and other metals would be recovered from the solutions by solvent extraction as they are circulated through a surface plant in a closed system. This provides for control of pH as well as containment of radioactivity. Radioactivity levels in the solution are low, and no radioactive impurities are expected in the copper. This technology will allow production from deposits too deep to be of interest with present technology and will reduce or eliminate some environmental effects of conventional mining, such as open pits, mine dumps, tailings ponds, and SO2 effluent. Close proximity of cities and towns is a limitation because of seismic damage. The technology requires further development for commercial application, and exploration targets must be enlarged to include deeper deposits, before the technology can be applied on a large scale.

Jan 1, 1974

By F. L. Ransome

THAT geology and mining should stand in close and mutually helpful relationship was recognized in the United States at an early period, but it is not possible to set a precise date at which the study of ore deposits became a thoroughly reputable and important branch of geological science. As early as 1850, appeared the first ,part of Foster and Whitney's report1 on the Lake Superior region, devoted mainly to the copper-bearing areas, followed, in the succeeding year, by the second part,2 on the iron lands. This appears to have been the first comprehensive report on what might, at that time at least, be called "western" ore deposits to be prepared and published under the auspices of the Federal government. An outgrowth from this work was Whitney's Metallic Wealth of the United States,3 which appeared in 1854. It was the earliest general work on the ore deposits of the country, and remained a standard reference book for marly forty years. From 1853 on, various parties were sent out by the War Department to explore routes for proposed railways and wagon roads from the Mississippi Valley to the Pacific Coast. Geologists were attached to most of these parties, but their work was mainly of broad reconnaissance character, little attention being paid to ore deposits. Although the

Jan 1, 1933

By K. E. Barrett

Exxon owns a uranium mill and holds two mining leases in Live Oak County, Texas, about halfway between San Antonio and Corpus Christi. The properties make up the Felder Uranium Operations which was reopened earlier this year. Exxon held an oil, gas, and other minerals lease on the J. C. Felder tract, which was adjacent to a relatively shallow uranium discovery by Susquehanna-Western, Inc. on the Marrs-McLean lease immediately south of the Felder property. Drilling in 1967 and 1968 confirmed the presence of reduced uranium mineralization in the basal sand unit of the Oakville formation on the Felder tract, which formed the major part of the roll-front deposit. In 1969 Exxon and Susquehanna-Western, Inc. entered into a sale and purchase agreement which provided for Susquehanna to mine and process Felder ore and purchase recovered uranium. Susquehanna moved an alkaline-leach mill from Wyoming, erected it on the Ray Point property, and placed it into operation late in 1970. Susquehanna mined and processed ore from the Felder and McLean properties through March 1973. Susquehanna ceased operations in March 1973. Exxon then acquired the mill and mill property. Exxon also purchased the mineral rights to the McLean lease, re-negotiated a mining lease for that property, and carried out shut-down programs for the mining and mill areas in the fall of 1973. The project was put on a standby basis until late 1973, when Exxon initiated mine feasibility studies for the project. MINE PLANNING EVALUATION The feasibility study for reopening the Felder Project began in late 1975 and was not completed until late 1976. I will discuss several areas of the feasibility study that required additional work prior to making the decision to renew operations. Ore Reserves Preparations for estimating the ore reserves began with the re-evaluation of more than 1500 natural radioactivity logs from exploration and pre-development drilling that had been completed on the property. These gamma ray logs of non-core rotary drill holes were the principal source of data used in making the estimate. Chemical assays of cores from the deposit were also used in the reserve determination. Electric resistivity and self-potential logs were made along with the gamma ray log. In December 1975 an additional core drilling project was undertaken to confirm the in-place density and radiometric equilibrium characteristics of the ore deposits. Comparison of chemical assays of cores with the U308 values calculated from the logs showed that the unoxidized ores were in radiometric equilibrium. In contrast, cores from anomalies occurring in near surface, weathered, and oxidized zones were in radiometric disequilibrium. Several important decisions were made in developing a mine plan or schedule of production from the Felder and McLean ore bodies. Disposal of Produced Mine Water: The ore bodies of the Felder Uranium Project occur at a point below the ground water table. The ore zones to be mined must first be dewatered to allow removal of mineralized material. In the open pit operations, this is accomplished by maintaining a perimeter ditch around the periphery of the open pit, allowing the interior of the pit to drain and collect into a sump and be pumped from the mine. In addition to anticipated water production from future mining operations, approximately 200M gallons of water was contained in three open pits left from prior mining operations. In two of these existing pits, the water was to be removed and disposed to allow for planned backfilling of waste material into these pits. The third pit would also have to be drained to allow continued mining of an area left from the prior operations. Essentially no ground water information was available for this area. The only data available was water production history from Susquehanna's mining operation. Two water wells were drilled early in 1976 on the Felder lease for use in obtaining hydrological data. A long term draw-down test was performed by pumping one water well and measuring water level drawdown in both the pumped well and the observation well. From these data, values for permeability and storage coefficient were calculated. These data were then used in modeling the aquifer to allow calculation of water influx into the mining area versus time. Once a schedule of water production, including the stored volume in the existing pits was calculated, alternate solutions for disposal were evaluated. The first system evaluated was a series of deep injection wells. The wells were designed to inject at a depth of approximately 3500 feet. Again very little information concerning reservoir characteristics of the receiving sand units was known. Using assumed values for reservoir permeability and storage coefficients, an injection well system was designed to allow for disposal of produced mine water. The biggest

Jan 1, 1979

By G. L. Corrigan, J. R. Coleman

Four factors that largely determine the end product obtained when cement and water are mixed are the chemical compo ition of the cement, the fineness to which the cement is ground, the amount of mixing water used, and the procedure used in mixing the slurry. In order to study the effect of two of these factors, the other two, the chemical composition and method of mixing, are kept constant, leaving as variables fineness and water-cement ratio. Through years of experimentation and usage it has been well established that the reduction of cement clinker to a very fine state by grinding improves its cementing value, principally because the reaction between cement and water takes place only at the surface of the cement particle, further action being hindered by the accumulation of reaction products that tend to coat the unreacted material. Therefore, the more finely ground a cement, and the greater the surface exposed in proportion to its mass, the more rapid is the hydration rate. Similarly, the greater the surface exposed, the greater the proportion of the cement that reacts, and the lower the amount of residue that is unreacted and so remains inert and unable to make any contribution to cementing properties.1 Although it is more expensive to manufacture a finely ground cement than a coarser cement, the advantages of fine grinding of construction cements were so marked that the fineness of cement has increased considerably in past years. When the manufacture of special oil-well cements was begun, it was necessary to consider properties of cement that previously were considered of minor importance, foremost of these being pumpability. Since, with the water-cement ratio constant, a finely ground cement will produce a neat slurry of higher consistency and shorter time of pumpability than will a coarser cement, there is a tendency to grind oil-well cements to a lower degree of fineness. Therefore, in order to produce oil-well cements that will make a slurry having a long time of pumpability, cements are manufactured that because of their coarseness generally are poorer cementing materials than some of the more finely ground construction cements. On the other hand, often it is necessary to pump oil-well cements for long periods of time to place them properly behind a string of casing in a deep well, and since high strength is not of foremost importance for such cement, there is room for considerable controversy concerning the proper fineness range.

Jan 1, 1941

By B. D. Crawford, S. A. Falconer

ONE of the outstanding achievements in connection with this country's war efforts has been the ability of our mining industry to supply from domestic sources many of the minerals of strategic importance that formerly were obtained principally from abroad. In order to meet the war-born demands for these scarce minerals, it has frequently been necessary to develop special methods for treating ores and waste products hitherto considered too low-grade to be workable; and to improve the efficiency of older beneficiation processes. Important in this respect have been the recent technological advances in the field of froth flotation of nonsulphide minerals. Because of the refractory nature of many of these domestic ores, and, in many cases, the limited available technical data regarding their treatment, much research work has been necessary. The following examples describe the procedures and reagents used, and the results obtained, in certain laboratory flotation investigations conducted by American Cyanamid Company's engineers at the Cyanamid Ore Dressing Laboratory on ores containing such minerals as scheelite, manganese oxides and carbonate, ilmenite, brookite, kyanite, garnet, spodumene, mica, feldspar, celestite, fluorite, chromite, corundum, and cassiterite. In presenting these data, we wish to point out that many of the minerals listed have previously been tested, and their amenability to flotation reported, by other investigators-notably those of the United States Bureau of Mines. In some instances the procedures and reagents used in the treatment of some of the following described ores follow rather well-established lines, and with one or two exceptions, the reagents used were well known. However, many of the ores tested did not respond well to simple conventional procedures and it was necessary to adopt special methods of conditioning and desliming, reversal of the usual order of floatability of some of the minerals, multiple cleaning, etc., in order to secure separations. Except where otherwise noted, all numbered and trade-marked reagents mentioned in this paper are marketed by the American Cyanamid Co. or the American Cyanamid and Chemical Corporation. TUNGSTEN ORES Tungsten minerals, vitally important in the war industries of the United States, rank high in the list of those we formerly had to import in order to supply our needs. Most of our domestic ores of tungsten contain scheelite as the valuable mineral. Today there are in the neighborhood of twelve tungsten operations in this country where flotation is used as the main or accessory form of benefication. It has, of course, been known for some time that scheelite and other tungsten minerals respond to treatment by flotation.

Jan 1, 1944

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400' to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400' to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941

THE fall meeting of the Petroleum Division at Fort Worth, Texas, Oct. 19 and 20, attracted even a larger number of members and guests than had been expected. Over three hundred registered the first day though the attendance and interest in the program was so keen that it was necessary to begin technical sessions without. waiting for registration. Members were present from as far away as Venezuela and while the larger attendance was from the Mid-Continent and Texas fields, California and the eastern states were represented by influential delegations. One pleasing feature of the meeting was the presence of a number of the older executives of the oil industry who evi-dently have come to realize the importance of the work the Division is doing in advancing technology. The meeting was held under the auspices of the Production Engineering section of the Division. E. 0. Bennett, chief production engineer for the Marland Oil Co. of Texas, and chairman of the Fort Worth Sec-tion. of the A.I.M.E., served as general chairman. J. B. Umpleby, president of the Petroleum Reclamation Co. and vice-chairman of the. Division was in charge of. the program with a special committee in charge of the technical sessions, of which A. C. Rubel of the Union Oil Co. was chairman. The sessions were held in the ball room of the Texas Hotel, which also served as general headquarters. The Wednesday morning session was devoted to the air-gas lift, John M. Lovejoy, chairman of the Division, presiding. The first paper, on the "Principles Under-lying the Air-Gas Lift." by J. 0. Lewis, was presented by R. R. Brandenthaler and discussed by Messrs. Love-joy and Millikan. Mr. Brandenthaler's own paper on the "Effect of Gas Lift on Physical Properties of Oil" followed and was discussed by Messrs. Umpleby, Walker, Mills, Bennett, Williams and Van der Gracht, the dis-cussion turning mainly on what is known and what unknown regarding oxidation of oil under various con-ditions. The third paper of the morning, "Handling Recirculated Gas," by R. D. Gibbs, was presented by Mr. Taylor. The discussion that followed was general and brought out a large amount of data on the costs and results of compression versus, absorption in total gasoline recovery from gas. The next group of papers dealt with "Mechanical Installations of Air-Gas Lift," the first by E. V. Foran on the Texas district outside the Gulf Coast, the second by L. L. Brundred on the Gulf Coast, the third on Kansas and Oklahoma other than Seminole by R. W. Bond, the fourth on Seminole by C. R. Swartz, and the fifth relating to California by H. C. Miller. Some of these were read by title in order to give more time for discussion which was active and participated in by. Mills, Griswold and others.

Jan 11, 1927

By John C. McDonald

This report is a continuation of an earlier one with a similar titlell to which the reader is referred for such details of procedure as do not appear here. A brief summary will be given of the objects and methods of the research at this time. It was pointed out in the previous paper that properties of magnesium alloys, as reported by different workers, often differ. This was thought to be due to differences in the preparation of the specimens tested. The ideal way to compare alloys would be to prepare each of them in so many ways that reasonable assurance would exist that the best properties of each alloy had been developed. The alloys would then be compared on the basis of these properties. The treatment actually given the alloys of these studies is necessarily a compromise with completeness. Nevertheless, enough treatments have been given so that a considerable degree of assurance exists that the properties obtained are representative of the alloy as such, and not merely of the manner in which it was worked and heat-treated. Procedure The alloys were cast in the form of cylindrical ingots, which were extruded into bar before being rolled. Some of the ingots were made in 2-in. diameter molds whereas all of the alloys of the previous paper were cast in 4.in. molds. For the larger ingots, the extruded bar was 2 by 0.140 in.; lor the 2-in. ingots, it was I by 0.100 in. Preliminary work indicated that strip rolled from both sizes of the bars would have substantially the same properties. The bar was rolled at some temperature in the range 400' to 800°F., for a total reduction of 50 per cent. The temperature and reduction per pass were so selected that no cold cracking occurred, but were such as to leave the metal in a semihard condition. The hot-rolled metal was annealed at several temperatures in the range 400' to 800°F.; it was also cold-rolled, and given solution and aging treatments. Two standard test bars corresponding to each treatment were pulled in tension, and the average was obtained. Twelve treatments, in all, were given each alloy. The properties of the annealed specimen having the highest percentage of elongation were selected to represent the alloy in the annealed condition. The properties of the cold-rolled specimen having the highest yield strength were selected to represent the alloy in the cold-rolled condition. These properties were then plotted against percentage composition by analysis. Smooth curves were drawn among the points, thus graphically averaging the results over the entire range studied. The percentage of elongation in the cold-rolled state was not plotted, since it was always low, and varied in an irregular manner with composition. Alloys Studied Magnesium-calcium Alloys.—Fig. I presents the results on the alloys of the rnag-

Jan 1, 1941

By E. W. Dewing

The solubility of A14C3 in cryolite melts has been determined by holding the melt with aluminum in a graphite crucible for 4 hr at temperature, quenching the crucible in water, and analyzing the frozen salt by CH4 evolution. The solubility is given by log(wt % Al4C3) = 1.298 -1.l74 (wt ratio NaF/AlF3) independent of temperature in the range 985" to 1080°C. Additions ofA1203, .NaCl, CaF2, or LiF up to 10 wt pct do not affect the solubility, provided that LiF and CaF2 are replaced by equivalent NaFin calculating the ratio. The variation with ratio can be accounted for quantitatively on the hypothesis that the Mechanism of solution is Al4C3 + 2AIF3 = 3Al2CF2 since the solubility varies as (aAIF3) 2/3; Mechanisms involving C", C2- , AlC- , A12C2+, and Al2CF+are re-jected. At ratios above 1.5 truces of c;- are also present. It was noticed some years ago in this laboratory that carbon specimens made cathodic in a cryolite bath decrease in diameter over a period of time.' It was also found that cryolite which had been melted with aluminum in a graphite crucible was sometimes yellow, and that A14C3 could be seen intimately embedded in the cryolite crystals. CH, could be recovered by acid attack. There was thus reason to believe that A14C3 had a genuine solubility in molten cryolite, and the object of this work was to measure it. EXPERIMENTAL AND RESULTS A graphite crucible, 2 in. ID and 2.5 in. deep with a close-fitting lid, contained about 150 g of cryolite (plus additives) and about 20 g of 99.99 pct pure Al. It was placed in a vertical Inconel furnace tube, covered with refractory brick to provide thermal insulation, and the top of the tube was roughly sealed with disks of asbestos paper. An alumina tube led argon down to the level of the top of the crucible to expel air, and an alumina sheath with a Pt-10 pct Rh ther-mocouple inside was positioned between the crucible and the wall of the Inconel tube. The tube was heated in an electrical resistance furnace for the desired length of time; the refractory bricks were then removed, and the crucible was lifted out with tongs and quenched in water. After such a crucible was broken open it could be seen that its inner surface was yellow owing to the presence of A1,C3, and sometimes needle-like crystals could be observed (at X75 magnification) growing out into the nielt. It was concluded that Al,C, had been present as a separate phase when the crucible was at tenlperature and that the melt had been saturated: in

Jan 1, 1970

By Charles H. White

Discussion of the paper of A. L. Sweetser, Tress., xxxviii., 236. CHARLES H. WHITE, Harvard University, Cambridge, Mass. (communication to the Secretary*) :-Those interested in leaching-processes gladly welcome contributions on either laboratory-tests, such as are reported by Mr. Sweetser, or mill-tests made on larger quantities of ore; and especially valuable are such reports, even with negative results, if the methods adopted are the outgrowth of an intimate knowledge of the conditions to be met. It seems very desirable, therefore, that as complete details as possible be given in reporting on such tests, particularly with respect to the character of the ore under examination. The questions of prime importance in the sampling of ore and in the making up of charges for laboratory-tests are those respecting its grade and uniformity. In Mr. Sweetser's paper the first of these questions is answered definitely, but the answer to the second, although apparent on inspection of the results, might be more strongly emphasized with advantage, since it is only on account of the extreme uniformity of the ore that the satisfactory results obtained could have been expected. It seems advisable to call attention to the fact that for average ores, and more particularly for uneven, or 11 spotted ores, the quantities used by Mr. Sweetser are not usually regarded as sufficient to insure satisfactory results. Let us examine the method somewhat in detail. The ore was crushed to pass a 5-mesh screen, and a sample of 45 lb. was taken out for assay and tests. The diameter of the largest grain of this ore, according to Richards,1 was 2.67 mm.-assuming here, as in the following cases, that the wire of the screen was of the largest size used for the purpose, thereby giving any possible advantage in this respect in favor of the method. The table in Richards's book, p. 852, shows that 45 lb. is a sufficient quantity for a fair sample with uniform ores, but not sufficient

Jan 7, 1908

By W. L. Zeigler

The mill is a departure from gravity concentration and has gained a reputation for the low initial cost of erection, extreme simplicity and the low cost of milling on the refractory character of the ore which is treated. The structure is of heavy timbers with board walls and corrugated roofing. The floors are concrete and the general layout is of the ground floor only. All the machines are driven by individual motors and the transmission usually found in such a plant is entirely absent. (Fig. 2.) ORE The ore from the Galena mine is sheared Wallace quartzite replaced with quartz, siderite, pyrite, sphalerite, gray copper (tetrahedrite) and very fine-grained argentiferous galena. Much of the silver is undoubtedly associated with the gray copper, the amount, however, apparently varies with different localities in the vein. An average assay of the ore is 6.0 per cent. lead, 1.3 per cent. zinc, 0.3 per cent. copper, 17.5 per cent. iron and 4.5 oz. silver. The concentrates produced are of an average 53 per cent. lead, 40 oz. silver, 8 per cent. zinc and 11 per cent. iron, while the tailings are 0.5 per cent. lead, 0.4 oz. silver. With the present market for zinc no attempt is made to produce a zinc concentrate.

Jan 1, 1927

By Eifu Moriyama

In 1953, Akita Plant, having a capacity of 560 tons per month, was built at Akita city by Mitsubishi Metal Mining Co., Ltd. for the purpose of treatment of the Ikuno and Akenobe Mines's zinc concentrates. Akita city, in whose neighborhood the black ore has been discovered, is located about 500 Km north of Tokyo and faced with the Japan Sea. This plant has been continuously expanded to have the capacity of 7,500 tons per month by the end of 1969. Efforts have been made to mechanize installations and to achieve continuous process. One of the characteristics of Akita Plant is the recent establishment of the residue treatment process. Many processes had been investigated, and finally commercial sulphate roasting of the residue has been carried out since 1965. At Akita Plant, special high grade electrolytic zinc, zinc alloy for die-casting (MAK brand), cadmium, cement copper and sulphur dioxide gas are produced. Sulphur dioxide gas is sent to the next door fertilizer plant to make acid. Table 1 shows production and the average assay of the electrolytic zinc and cadmium in 1969. This plant is divided into four main sections, two subsections, and residue treatment sections: Main sections are: (M-1) roasting section, (M-2) leaching and purification section, (M-3) electrolysis section, (M-4) melting and casting section,

Jan 1, 1970

By David L. Riley

Organization The Foundling Years - 1901-1921 Two events that occurred in the decade of the 1870's have had far- reaching effects upon the petroleum industry in the U. S. and formed a basis for the birth and growth of the petroleum engineering profession. In 187 1 a group of mining engineers banded together to form the American Institute of Mining Engineers "to promote the arts and sciences connected with the economic production of the useful minerals and metals and the welfare of those employed in these industries by all lawful means". Although it was to be 30 years before an event in Jefferson County, Tex., would cause petroleum to come into prominence as the world's principal source of energy, the founding of AIME provided an organization through which the application of engineering to the recovery of petroleum and natural gas could be developed. In 1875, a Pennsylvania judge who was called upon to make a decision for which there was no precedent reasoned that oil and gas were like wild game - they belong to the man who can reduce them to possession. His decision sired the "rule of capture", which stated that a landowner possesses the oil produced from his well, even though it might have come from beneath his neighbor's land. This interpretation prevailed for more than half a century, during which time gold-rush tactics were employed in the drilling and production of many oil fields in the U. S. The resulting overproduction and tremendous waste led to a virtual economic collapse of the U. S. petroleum industry by the early 1930's. To bring these malpractices under orderly control, proration and conservation policies were developed, and petroleum engineering became the medium through which they were devised and implemented. At the turn of the 20th Century, more than 90 percent of this country's oil production came from the East, but the picture changed with dramatic suddenness on Jan. 10, 1901. Captain Anthony F. Lucas, an Austrian mining engineer, had noticed concentrations of oil and gas near salt domes during mining operations on the Louisiana Gulf Coast. In 1899 he leased a low mound near Beaumont, Tex., began drilling, and after almost 2 years of intermittent operations, brought in the famous "Lucas Gusher"

Jan 1, 1971

By Dr. O’Neil Thomas J., Donald W. Gentry

"There is nothing so disastrous as a rational investment policy in an irrational world. " John Maynard Keynes INTRODUCTION The principles of time value of money concepts were discussed and illustrated in Chapter 3. The importance of time value of money and the role it plays in the discipline of engineering economy have been well established. Chapter 9 further illustrates that project evaluation criteria that fail to consider this aspect are seriously deficient in providing appropriate capital budgeting decisions. Obviously future project receipts and expenditures must be discounted to permit valid comparisons with current cash flows. Although the concept of discounting is widely accepted, the selection of the appropriate discount rate has been the source of considerable debate and much disagreement. This chapter, then, provides an overview of the theoretical and practical considerations for determining or calculating the appropriate rate for discounting, the corporate cost of capital. Note specifically that the assumption of constant business risk among the various investment alternatives is maintained throughout this chapter. THE INTUITIVE APPROACH The fact that interest exists suggests that all money has a cost associated with its use. The cost of this money may be the result of either explicit or implicit charges or some combination of the two. Any time an individual or a firm consumes less than the total earnings generated in a given time period that individual or corporation has either consciously or unconsciously decided to "invest" these excess funds in some type of activity. For instance, the individual or corporation may decide to invest the excess funds in tangible property, place the funds in a money market account, or simply keep these funds in cash. In any case, these funds have been invested in some type of activity which promises the investor some return on the invested capital. Remembering that the firm's primary objective should be the maximization of shareholder wealth, then intuitively it seems reasonable to suggest that the firm should not invest in any project where the anticipated return does not exceed the cost of funds (capital) committed to the project. Indeed, if the firm always invests in projects having returns in excess of the cost of capital committed to them, then the wealth of the firm (as measured by the price of common stock) should be in-

Jan 1, 1984

By E. J. Harris

Mathematical solutions to complex mine ventilation problems are possible, but often the airway network is so complex that the mathematical solution becomes tedious and impractical. A fluid network analyzer, designed and built for analyzing mine ventilation problems has been in service at the Bureau of Mines' Pittsburgh station for approximately nine years. Using this analogue as a model, the mine ventilation network is simulated electrically by a combination of series and parallel circruits laid out to conform with actual mine airways. A tungsten filament lamp, referred to as a Fluistor, is used to simulate mine airway resistance. With the analogue ventilation model established, voltages of the proper amplitude to represent ventilating pressures are impressed across the circuit at points where mine fans and airshafts are located. Comparison of mathematical solutions of complex systems against analogue results showed a maximum variance of 3% for pressures and 2% for quantities. An electrical fluid network analyzer, designed and built especially for analyzing mine ventilation distribution problems, has been in service at the Bureau of Mines, Pittsburgh, for approximately nine years. It is a nonlinear, low voltage, fluid network analyzer of the type developed by the late Malcolm S. McIlroy, Professor of Electrical Engineering, Cornell University, who cooperated with G. E. McElroy, of the Bureau of Mines, in adapting the instrument to mine ventilation systems. Several modifications have been made since the original installation, but a considerable part of this paper is drawn from G. E. McElroy's original description of the analogue.' The analysis of water or gas distribution systems2,3 led to the development of this type of network analyzer. Several similar units are now employed by utility companies for this purpose; however, the Bureau of Mines unit is the only one designed specifically for mine airflow problems. Other airflow analogues employing a similar principle have been used at the Central Research Station of the Netherlands State Mines,4 in England,5 and in South Africa.6 Information on these devices indicated they were somewhat inflexible because commercial lamps with a sufficient range of resistance are difficult to obtain. Computers for mine ventilation analysis have been developed in Germany which instead of lamps use a variable resistance to adjust for turbulent flow laws. These units are expensive, but excellent results have been reported in their application. THE ANALYZER Theory of Application: As airflow generally follows the law of turbulent fluid flow, resistance to flow is nonlinear; consequently, the problem has been to find a nonlinear resistance element of suitable range that can be used for electrically simulated airflow. Tungsten filament lamps operated on alternating or direct current approximate the square-law resistance characteristics of mine airflow over a large range below maximum or rated voltage; that is, the voltage drop varies approximately as the square of the current. Consequently, the heart of the network analyzer is a nonlinear resistor known as a Fluistor, which is simply a custom-made low voltage, tungsten filament lamp that is available in a progressive series of relative resistance values ranging from 0.05 to 500 in nominal 5% steps. However, variations in manufacturing large groups result in differences of 1 to 3%, but series arrangements required for high-loss branches can be matched within about 1%. Utilizing a combination of Fluistors and load circuits, the mine ventilation system is duplicated electrically. Intake load circuits are connected from power intake to primary points of the circuit network; segments of unregulated flow along intakes and returns are represented by Fluistors, regulated splits and leakage paths are represented by load circuits and Fluistors of proper capacity; and mine exhausts are connected to ground from the last point of the network to complete the circuit. For the special purpose of representing booster fans or natural draft conditions, boosters or separate source circuits are provided that can be inserted between any two points of a network to increase voltage to the required value. Physical Layout: The analyzer consists of three 42

Jan 1, 1963

By Harry C. Swanson

TRANSPORTING concrete from mixer to forms has always been a problem. Twenty-five years ago this task was generally accomplished by means of wheelbarrow or concrete buggy. On large dam jobs, as the number of these projects increased, the gantry crane or highline came into use. Today several methods of handling concrete are employed on smaller surface construction jobs, for example, transit-mix trucks or dumpcrete trucks, which have crawler cranes with buckets for placing concrete into forms. In 1944, during early stages of developing Mather mine A shaft, several large underground concrete jobs were necessary. At this time the Cleveland-Cliffs Iron Co, purchased the first pump-crete machine, introduced by the Chain Belt Co. of Milwaukee. The machine was used to pour approximately 200 cu yd of concrete for a dam, or bulkhead, located 400 ft from the shaft. Concrete was mixed on surface, lowered down the shaft 1000 ft in a 2-cu yd bucket hung under one skip, spouted into the bowl of the pumpcrete machine from the bucket, and pumped directly into the forms. Since the day of the first pipeline concrete in 1944 to the present time, other equipment and other methods have been developed to permit transportation of concrete by pipeline through vertical and horizontal distances totaling 1 mile from mixer to forms. Much of the efficiency in present handling of underground concrete can be credited to the Bethlehem Cornwall mines, where concrete was transported through 6-in. pipe for great distances down an inclined shaft and along levels into forms.' During initial development of Mather mine B shaft, with concrete work under way on two or more levels at one time, the pneumatic concrete placer, Fig. 1, was selected as best adapted for underground concrete transportation. The 3/4-cu yd pneumatic placer is a small machine readily moved from one location in the mine to another. It can be equipped with two sets of mine car wheels, which will permit moving on regular mine tracks. It is therefore possible to send concrete through the pipe at great velocity; the pipeline is clean after each shot except for the film of cement adhering to the inside. With the proper slump in the concrete, it is possible to shoot concrete 2000 ft with this machine, using the mine supply of compressed air at 95 psi. This equipment was first used at Mather mine B shaft to concrete slusher drifts, Figs. 2 and 3, and finger raises located about 2000 ft from the shaft. In several instances there were bends into crosscuts and up vertical distances into the forms. For the first pours two placers were used. The first was located near the shaft where the concrete could be spouted into it from a 2-cu yd concrete bucket on the cage. The second was set on the side of the drift at a point approximately 1500 ft from the shaft. The concrete was shot directly into the second placer from the first unit and from the second machine directly into the forms. After completion of several pours with the two machines, a trial pour with only one placer located at the shaft proved that the second placer could be eliminated. Since then all pours have been successfully completed with only one placer underground. As production of iron ore from the mine increased and the development program expanded, use of the cage for handling mine supplies and concrete became a major problem. This brought about the first attempt at shooting concrete vertically down the shaft for 2600 ft. A 6-in. pipeline with victaulic couplings installed during shaft sinking was used for the trial. One placer was set on surface 250 ft from the collar of the shaft so concrete could be spouted directly into it from the mixer. This machine shot the concrete 250 ft horizontally on surface to the shaft, 2600 ft vertically down the shaft, and 100 ft horizontally into the second placer located near the rib of the shaft station or plat. The second machine shot the batch into the forms, about 2000 ft. Total distance horizontally and vertically was 4800 ft. The entire time cycle for a ¾-cu yd batch of concrete from the mixer on surface to the forms underground totaled about 5 min. During the past two years the two-placer method from the mixer on surface to the forms underground has proved a very efficient means of transporting underground concrete. Advantages of using pipeline concrete are as follows: 1—Interference with normal mining operation is eliminated. When the cage, skips, mine cars, or mine openings are used for transporting concrete and materials used for making concrete, mine operation suffers in one way or another.

Jan 1, 1955

By Fred C. Bond, Walter Maxson

Since the last publication of tabulated results of grindability tests by the authors1 the total number of ball-mill tests made has more than doubled, and rod-mill tests have become increasingly important. Nearly all of the standard closed-circuit ball-mill and rod-mill tests made to date are included in the present tabulation, together with additional information not previously published. Standard Ball-mill Grindability Tests The method of conducting these closed-circuit tests has not been altered. A sample of the ore or other material is stage-crushed in rolls set at 1/16-in. opening with a 6-mesh screen until all of it has passed the screen. The combined minus 6-mesh screen undersize is mixed, sampled and screen-analyzed, and its apparent specific gravity is determined by packing and shaking in a standard container, and weighing. The apparent specific gravity ordinarily is about 60 per cent of the true specific gravity. The unit volume present in the mill .in all tests is 700 C.C. of the packed minus 6-mesh roll product, and the number of grams occupying 700 C.C. is the unit test weight. This weight is placed in the mill dry, ground for the number of revolutions estimated to be necessary, discharged, and screened mechanically in three testing sieves of the mesh size at which the test is to be conducted. The oversize, or circulating load, is weighed, sufficient fresh minus 6-mesh feed is added to bring the total weight up to that of 700 c.c., and the charge is returned to the mill for a second grinding period. All standard ball-mill tests are conducted at 250 per cent circulating load, and the number of revolutions in the mill necessary to obtain this circulating load at any grinding period is estimated from the results of the previous period. The number of net grams of screen undersize produced per revolution of the mill approaches its final equilibrium value after several grinding periods, and this is recorded as the grindability, or relative ease of grinding, of the sample. A cylindrical ball mill, 12 by 12 in. inside, with a smooth lining and rounded corners, is used with a revolution counter Under standard grinding conditions it runs at 70 r.p.m., and contains a charge of 285 iron balls weighing 20,125 grams, ranging from I½ to ¾ in. in diameter. Tests are conducted at the mesh size to which the ore is to be ground in practice, and the capacity- to be expected from a given mill is calculated by comparing the grindability with that of a similar standard ore whose performance characteristics arc known. No satisfactory general formula for convertillg standard grindabilities into corn; rnercial power-consumption values has been

Jan 1, 1943

By Frank G. Norris, William C. Marshall

The term "total operation" is meant to include problems that cannot be answered from the standpoint of either the blast furnace or the open hearth separately but must be studied by considering the interrelations of these two processes. There are other ways of expressing this idea, such as reference to the blast furnace and the open hearth as an integrated process or by emphasizing the interrelation and interaction of the two processes. Much progress has been made in metallurgy and other fields by breaking complicated processes and reactions into simpler parts and studying each of these intensively. This intensive study frequently is done under the controlled conditions of the laboratory experiment. The entire process is understood better by this specialized study of each of its parts. This procedure is so satisfactory that an understanding of unit operations is a fundamental requirement of chemical engineering. Much discussion of the open-hearth process is in terms of one heat or one furnace. By this is meant that the conclusions apply to one furnace, wherever it may be. Furnaces seldom occur singly; they are located in groups and these groups are located in steel plants having characteristic blast-furnace equipment. Problems that are affected by the number of furnaces operating, by the number of blast furnaces in the plant and by the fact that the open- hearth shop is dependent on the bla furnace capacity rather than on its o needs for the amount of iron available among those which, in contrast to "u operations," we have chosen to call "to operations." Many of these problems, such as cha ing delays and the general mechani problem of material handling, are familiar that merely to define " total op ations" brings them to mind. These pr lems are of such importance that alrea they have received careful thought 2 study and probably will merit furt study. As will be shown later, these mechan problems are so intimately related metallurgical problems of total operati that the effect of certain changes cannot ascribed with certainty to either a mech ical or a metallurgical cause entirely. Although the problems coming under general term of "railroading" are inclu in the problems of total operations, ' others are selected for the discussion of paper, One of these is strictly a problem composition; the other is strictly a prob of tonnage production. The problem of phosphorus is pres whenever basic open-hearth slag is used the blast-furnace burden. This practice not secured universal acceptance chi because of difference in preference for h manganese or low-manganese pig iron hjgh-manganese (2.00 per cent) iron is to made, open-hearth slag as a source manganese is usually cheaper than m

Jan 1, 1943